Oxygen smelting of copper or nickel sulfides

ABSTRACT

Copper and or nickel sulfide ore concentrates of high intrinsic value are oxygen smelted by introducing such concentrates via feeds (32) into a closed loop extraction circuit in which a molten sulphide carrier composition is forcibly circulated by an R-H unit (14) through lower hearth (10) at which feed of the concentrates takes place, and upper hearth (12) at which controlled oxidation with technically pure oxygen through top lances (4) takes place to oxidize copper sulphide and iron sulphide in the concentrate to iron oxide. Copper is removed via line (42). Slag containing iron oxide is discharged via weir (22) into a secondary circuit in hearths (24 and 28). Carbonaceous reductant added to the slag in (28) reduces the iron oxide in the slag to hot iron product.

This invention relates to the oxygen smelting of copper sulphide oreconcentrates, nickel sulphide ore concentrates or bulk copper and nickelsulphide ore concentrates, and is particularly concerned with the directoxygen smelting of those of such concentrates which have a highintrinsic energy value.

Existing copper smelting technologies include:

Top Blown Rotary Converting

Electric Smelting

Noranda Continuous Smelting

Mitsubishi Continuous Smelting

Inco Flash Smelting

Outokumpu Flash Smelting

For those concentrates with relatively low intrinsic energy value (i.e.those where large net amounts of exothermic heat are not generated),high level oxygen enrichment can be used with the above technologies,and in certain cases depending both on the existence of highcopper/sulphur ratio and a low sulphide iron content in a borniticconcentrate, it is possible to produce blister copper directly in asingle step using technically pure oxygen ie oxygen of commercial purity(95% or higher).

Typical chalcopyrite copper concentrates, on the other hand, whenoxidised directly to blister copper, generate reaction heat in excess ofthat required for autogenous operation, even at low levels of oxygenenrichment. The inability to dissipate this additional heat with currentused technology prevents or complicates the use of higher levels ofoxygen enrichment.

Besides the excess heat generation problem already referred to, singlestep smelting of copper concentrate with relatively high levels ofcertain impurity elements (Arsenic, Antimony and Bismuth) tend toproduce final blister copper with unacceptable levels of theseimpurities because their elimination is generally lower than byconventional smelting routes. As a consequence direct smelting toblister copper has to date been restricted to relatively cleanconcentrates.

GB-A-2048309 (corresponding to EP-A-016595 and U.S. Pat. No. 4,334,918)discloses a process for recovering nonferrous metals from their sulphideores wherein a molten sulphide carrier composition (or matte) isforcibly circulated through an extraction circuit from which thenon-ferrous metal or its sulphide can be continuously extracted at anelevated temperature. The method involves introducing the sulphide oreinto the matte at an ore-receiving station so that the ore is dissolvedin or melted by the matte, and contacting the matte containing said orewith oxygen at an oxidation station so as to oxidise at least part ofthe ore and/or the matte, heat generated during the oxidation step beingrecovered by the matte and transmitted thereby to endothermic sites inthe circuit. GB-A-2048309 is not so concerned with copper sulphide oreconcentrates having a high intrinsic energy value, but it does disclosethe treatment of a copper-zinc ore concentrate containing 25.6% copper,10% zinc, 1.7% lead, 24% iron and 33% sulphur. With such an oreconcentrate, oxidising is effected in an oxidising unit divided intofirst and second parts wherein a major part of the matte passes throughthe first part and is oxidised by oxygen lances located above thecirculating stream, the oxidation being controlled so that onlypreferential oxidation of the ferrous sulphide occurs. A minor portionof the matte is directed through the second part where it is top blownwith oxygen-enriched air so that both Iron and copper. sulphides areoxidised to produce a molten copper phase as well as a slag phasecontaining iron oxides and some dissolved cuprous oxide. The moltencopper phase produced in the second part is separated so that part canbe extracted as blister copper and the remaining fed back to ade-zincing vessel. After passing through the second oxidising part, theremaining matte and slag phases are re-mixed in a cascade fashion withthe main matte stream in a slag cleaner, with coal being introduced intothe re-mixing region so as to reduce the oxygen potential of the slagand hence decrease the solubility of the cuprous oxide in the slag.Further slag cleaning is provided by addition of iron pyrites to theslag. It will be appreciated that, because such a copper-zinc oreconcentrate has a relatively high zinc content, it cannot be regarded asan ore concentrate having a high intrinsic energy value.

EP-A-0266975 (corresponding to U.S. Pat. No. 4,701,217) discloses ananalogous melt circulation process for reducing oxides of copper andnickel or copper, nickel and cobalt in deepsea manganese nodules usingcirculating copper nickel sulphide matte and a suitable carbonaceousreductant (eg partly devotatilized coal), and also discloses thepossibility of mixing such nodules with additional copper which may bein the form of a copper-bearing mineral such as chalcopyrite, or scrapcopper in order to ensure that phase separation of a nickel-copper ornickel-cobalt-copper alloy takes place.

It is an object of the present invention to provide an improved methodof direct oxygen smelting of copper sulphide ore concentrates, nickelsulphide ore concentrates or bulk copper and nickel sulphide oreconcentrates (hereinafter simply referred to as copper/nickel sulphideore concentrate) which can enable reduced Gas emissions to be achievedwhilst at the same time permitting effective use of the high intrinsicenergy of such ore concentrates.

According to the present invention, there is provided a method of oxygensmelting of a copper/nickel sulphide ore concentrate of high intrinsicenergy value, comprising the steps of forcibly circulating a moltensulphide carrier composition through a closed loop extraction circuitfrom which copper/nickel or sulphide(s) thereof can be continuouslyextracted at an elevated temperature, introducing the ore concentrateinto the molten carrier composition at an ore receiving station so thatthe ore is dissolved in or melted by the composition, contacting themolten carrier composition containing said ore with an oxidising gascontaining at least 30 vol % oxygen at an oxidation station so as tooxidise at least part of the ore and/or the molten carrier composition,and utilising heat generated during the oxidation step as a result ofoxidation of the ore concentrate.

Most preferably, the oxidising gas is technically pure oxygen, althoughan oxidising gas having a lower oxidising potential may be employed ifcircumstances permit.

in one embodiment, the generated heat is utilised by smelting thecopper/nickel sulphide ore concentrate of high intrinsic energy valuewith another mineral concentrate of low or negative intrinsic energyvalue, eg high-grade zinc concentrate, high-grade lead concentrate oreven a bulk flotation concentrate containing both lead and zinc,preferably low in gangue oxides for highest thermal efficiency, but notnecessarily restricted to high grade concentrates if metallurgicalefficiency dictates otherwise. With such a process, metallic copper,metallic zinc and metallic lead can all be obtained as products in theprimary smelting circuit employing forced circulation of copper/nickelsulphide through various extraction zones, following the teachings ofGB-A-2048309. The primary zinc and/or lead formation reactions consumethermal energy and so, if the ore concentrates are added in the correctproportions, the excess energy released on direct smelting ofcopper/nickel concentrate using technically pure oxygen can be balancedagainst the endothermic requirements of zinc and/or lead production.This has the advantage that the energy required for zinc and leadproduction is provided in situ within the smelter so that no externalfuel is required and all the benefits of virtually zero gas emissionsmelting are secured. Preferably, copper is extracted as the metal,whilst nickel is extracted as high grade nickel sulphide.

The use of technically pure oxygen or oxygen enriched air as opposed toair has the advantage that it is not necessary to treat large quantitiesof nitrogen to remove entrained pollutants such as oxides of nitrogen,oxides of sulphur and volatiles such as arsenic, antimony and bismuthbefore discharge.

It will be appreciated than the large amounts of oxides of sulphur whichresult from the process using technically pure oxygen are relativelyundiluted with other gases and can therefore be relatively easilyconverted to sulphuric acid.

Since copper/nickel sulphide ore concentrates also contain appreciablequantities of iron, the method of the invention produces relativelylarge quantities of slag as a major waste product. It is possible usingthe method of the present invention to reduce the quantity of slagsubstantially by reducing the iron oxide in the slag to form iron metal,thereby enabling the recovery of iron by utilising at least part of theexcess heat generated during oxidation of copper/nickel sulphide toprovide the necessary energy for the endothermic conversion of ironoxide to iron. This can conveniently be conducted by employing asecondary circuit utilising the molten slag phase removed from theprimary extraction circuit for copper/nickel or sulphide(s) thereof.

By using such a technique to recover iron as a marketable product fromthe slag, the mass of slag finally discarded as waste and its content ofdeleterious elements can be very substantially reduced and thus moreeasily meet increasingly stringent legislation on waste disposal.

In the secondary circuit, the molten slag is contacted with a suitablereductant, eg, coal, coke, lignite or natural gas, to produce a liquidiron product which contains some dissolved carbon, such liquid ironproduct accumulating below the slag on the hearth of the secondarycircuit. This liquid iron product can be tapped off intermittently orarranged for continuous removal via a siphon-type continuous tapper ofthe type used in blast furnace operations. By allowing the liquid ironproduct to accumulate as a continuous layer covering the entire furnacehearth, the deleterious effects of chemical attack by the normallyaggressive liquid slag and erosion of the refractory lining of thefurnaces is thereby prevented. Vertical walls and/or partitions employedin this secondary circuit can be conveniently. protected from erosionand chemical attack by freezing a protective slag lining on these areasby water cooling such walls and/or partitions. Because the slag melt iscontinuously circulating through the secondary circuit and overflowingback to the primary extraction circuit, the depth of melt can be kept toa minimum and thus the area exposed no water cooling is relatively smallas compared with the larger hearth areas protected with the hot metallayer, which does not require forced water cooling but may, ifnecessary, be cooled more moderately by forced air circulation ornatural convection.

As will be appreciated from the above, all the energy requirements foreffecting the chemical reduction of the slag to metallic iron arederived from the surplus exothermic heat released in the primaryextraction circuit. This surplus exothermic heat is picked up by theslag and transferred to the secondary circuit by forced circulation.Whilst in the secondary circuit, there is no requirement for a massivereduction in the iron oxide content of the slag as it is normallycirculated at a rate many times of that of the iron production rate sothat the overall temperature drop of the slag phase in passing throughthis secondary circuit is relatively small, typically not more than 50°C., although in certain cases there may be somewhat higher levels oftemperature change, depending principally on the level of temperaturerequired in the non-ferrous smelting circuit and also the liquidustemperature of both the slag and the molten iron/carbon alloy formed inthe secondary circuit. Because the change in iron oxide concentration isrelatively small, the steady state concentration of iron in the slag canbe maintained at a high level, thereby permitting favourable reductionkinetics to be established and thereby permitting high productivity.

Once the secondary circuit is established, no further siliceous or otherflux is normally required to be added. All that is happening is that theiron content of the slag is increased to a small extent in the ironslagging zone. By not having to add cold solid fluxing agentscontinuously, appreciable savings in process energy are achieved. As thegangue oxides introduced into the slag build up to a steady slate level,a small proportion is removed to bleed off non only these oxidesintroduced into the slag in the non-ferrous loop but also those derivedfrom the coal ash slag in the secondary circuit.

Because the iron in the slag is already principally in the ferrous stateby the time it is removed from the primary (extraction) circuit andbecause it is already at high temperature, the carbon requirements foriron production by this technique are considerably less than any otheriron-making technology. The carbon monoxide primary gaseous product,containing some carbon dioxide and coal volatiles (if coal is used), ismore than sufficient to fire a waste heat boiler by combustion with airof relatively moderate preheat whereby the steam so produced is enoughto generate the electricity requirements of air separation to oxygen andnitrogen. The oxygen is produced at a rate sufficient to supply therequirements of direct oxygen smelting of the copper/nickel sulphide andother mineral sulphides fed into the non-ferrous smelting circuit.

It is possible for so-called "tramp elements" to build up in the ironproduced because such tramp elements are carried over in the slag fromthe primary smelting circuit. By circulating the iron-containing slag asdescribed above, the build-up of tramp elements in the iron may bemitigated by restricting the amount of reductant added to the upstreampart of the secondary circulation circuit. In doing so, the hot metalinitially formed collects the residual copper and other reducibleelement impurities that leave the primary extraction circuit dissolvedin the slag. By collecting this initial small amount of hot metal andreturning it to the primary extraction circuit (probably intermittentlyby tapping off this "front-end" hot metal which is not allowed no mixwith the major portion of the hot metal), the non-ferrous metals sorecovered are usefully returned to the non-ferrous loop without creationof a waste or by-product stream. Following this initial reduction with arestricted amount of reductant, the slag melt is then contacted withfurther reductant to produce a low-impurity, high quality hot metal.This advantage cannot be achieved using conventional bath smeltingtechnology because the extensive back mixing involved in such processes.

Particularly in the case where the ore concentrate being used as feedcontains impurities such as arsenic and antimony, it is preferred tosubject such ore concentrate, possibly but not necessarily in crudepelletised form to facilitate solids charging and assimilation into thecirculating sulphide material, to direct thermal radiation from thecirculating matte/slag surface at high temperature before assimilationthereof into the matte. This radiant pre-heating of the feed solidsoccurs within the refractory enclosure and permits the so-called"labile" sulphur to be evolved from the solids under neutral or reducingconditions. The sulphur gas thus emitted takes with in arsenic andantimony sulphides which are released from the solid charge whilstundergoing pre-treatment. By the time the solids are assimilated intothe circulating matte, the level of impurities in the feed solids isreduced to such a low level that the final non-ferrous product (egblister copper) is well within the specification for these impurityelements. Because of the high thermodynamic copper activity in the meltof single step copper production processes, there is little scope forarsenic and antimony elimination from the matte once these elements haveentered it. Thus, the radiant pre-heating technique described aboveprevents arsenic and antimony from entering the matte and enables themto be eliminated as a gaseous species along with the elemental sulphurthe gas reduced from the reducing side of the circuit. If desired, thesegaseous species may be partially separated from the sulphur prior tocondensation.

In the accompanying drawings

FIG. 1 is a schematic view showing a method and apparatus for directoxygen smelting of a copper sulphide ore concentrate of high intrinsicenergy value to produce blister copper and to recover iron therein as aniron-carbon (liquid metal) product with low amounts of discard slag, and

FIG. 2 is a cross-section on the line A--A of FIG. 1.

Referring now to FIG. 1 of the drawings, a primary closed loopextraction circuit 10 for a molten carrier material (matte) of coppersulphide containing FeS and other metal sulphides is established throughfirst and second hearths 10 and 12 by means of an R-H degassing unit 14having an inlet snorkel 16 extending into the molten matte in a slagseparation zone 17 of the first hearth 10 and an outlet snorkel 18 whichsupplies molten matte into the second hearth 12. The slag separationzone 17 is connected with the remainder of the first hearth 10 by anoverflow weir 19 for matte and slag. Molten matte circulates back to thefirst hearth 10 from the second hearth 12 via a connecting passage 20fitted with means such as a gas (eg nitrogen) curtain 21 for preventingthe gases in the respective hearths 10 and 12 from intermixing.

The slag separation zone 17 has an overflow weir 22 for discharging slagfrom the first hearth 10 of the primary circuit to one end of a firsthearth 24 of a secondary smelting circuit. The weir 22 is fitted withmeans such as a gas (eg nitrogen) curtain 25 for preventing gas in thehearths 10 and 24 from intermixing. An R-H unit 26 serves to transferslag from the first hearth 24 into one end of a second hearth 28 of thesecondary smelting circuit. The opposite end of the second hearth 28 ishigher than the second hearth 12 of the primary circuit and dischargesinto the latter via weir 30 fitted with means such as gas (eg nitrogen)curtain 31 to prevent gases in the respective hearths 28 and 12 fromintermixing.

Pelletised copper sulphide ore concentrate C containing iron (egchalcopyrite) is passed via multiport feeds 32 into either side of thefirst hearth 10 of the primary smelting circuit. The ore concentrate Cmoves down sloping side walls 34 (see FIG. 2) of the first hearth 10 fora substantial distance as a relatively thin layer before entering thelayer of molten matte M which is circulating at high rate through thefirst hearth 10. During its passage down the sloping walls 34, thepellets of ore concentrate C are exposed to radiant heat within thefirst hearth 10 which, like the other hearths 12, 24 and 28, is renderedsubstantially gas tight by a refractory roof 36. Instead of usingsloping side walls 34, the ore concentrate C may be transported using aheat-resistant belt conveyor within the high temperature region of thehearth 10, which takes the concentrate C along the entire length of thehearth 10 before discharge into the circulating copper matte. The oreconcentrate C is thereby heated to a temperature of about 1000 K., atwhich temperature labile sulphur is volatilised as sulphur vapour andarsenic, antimony and bismuth present as impurities in the oreconcentrate C are also volatilised in the form of their sulphides. Suchvapours are removed from first hearth 10 via line 38 leading to acondenser (not shown). The ore concentrate 34, from which the arsenicetc have been volatilised, dissolves in and is melted by the circulatinglayer of molten matte M and passes over weir 19 together with slag andinto the slag separation zone 17. The matte is then transferred to thesecond hearth 12 through the R-H degassing unit 14. Since the inletsnorkel 16 extends below the layer of slag which has formed on top ofthe molten matte layer M, relatively clean matte is transferred to thesecond hearth 12 by the R-H unit 14. Slag overflows from zone 17 viaweir 25 into the first hearth 24 of the secondary smelting circuit.

The molten matte M transferred to the second hearth 12 of the primarysmelting circuit and which contains the dissolved ore concentrate C isthen subjected to oxidation using technically pure oxygen suppliedthrough top lances 40. Such oxidation is controlled so as to convert thecopper sulphide to copper metal and to convert the iron sulphide in theore concentrate C to ferrous oxide. The copper produced is bled off vialine 42 as blister copper whilst the matte layer masses back to thefirst hearth 10 via passage 20. Slag cleaning operation (eg using ironpyrites) is effected at this location in hearths 10 and 12. Sulphurdioxide produced during the oxidation step and any excess unreactedoxygen are passed via line 44 at the downstream end of hearth 12 to asulphuric acid production unit 46, the separated oxygen being recycledback to the lances 40 to be mixed with fresh technically pure oxygen. Arelatively slow-moving layer of molten copper is maintained in thehearths 10 and 12 below the rapidly circulating molten matte M so as toprotect the hearths against severe erosion which would otherwise occurif the circulating matte M were directly in contact with the hearth.

Whilst the above-described production of blister copper is taking place,slag containing large amounts of iron in ferrous form is circulatingthrough first hearth 24 and second hearth 28 of the secondary smeltingcircuit. At the upstream end of the first hearth 24, a controlledaddition of carbonaceous reductant is made through line 50 in an amountsuch as to reduce any copper oxides which have been carried over by theslag to copper. This separates out below the layer of slag and is bledback to the upstream end of second hearth 12 via bleed line 52. Toinhibit erosion of the hearths 24 and 28 of the secondary smeltingcircuit, a layer of liquid metal Fe (C) is maintained therein below theslag layer and walls and partitions of the hearths 24 and 28 are watercooled so as to cause a layer of solid slag to adhere thereto andprotect these parts against erosion and corrosion. The slag which hasbeen subjected to controlled reduction then passes along first hearth 24to be transferred via R-H unit 26 into the second hearth 28. Furthercarbonaceous reductant is added to the slag in hearth 28 via lines 58and 60 which are spaced apart longitudinally of the second hearth 28.The carbonaceous reductant introduced via lines 58 and 60 serves toreduce ferrous iron in the slag no iron metal containing carbon (hotmetal) which is bled off at the appropriate rate via line 62 as hotmetal product containing low amounts of impurities such as copper.However, if the copper level is unacceptably high, the hot metal may bede-coppered by sub-surface injection of sodium sulphate. This is reducedby the carbon in the melt to form a sodium sulphide slag which in turnextracts both copper and sulphur from the molten iron to form a slag ofCu₂ S dissolved in Na₂ S. After separating the slag from the refinedmolten iron, the sodium sulphate can be regenerated from the slag by airinjection, leaving two immiscible liquid layers, molten sodium sulphateand molten Cu₂ S, the latter being returned directly to the coppersmelting circuit. Discard slag is bled off via line 64 at the upstreamend of second hearth 28.

Since the slag in the secondary smelting circuit is circulated rapidly,the iron content of the slag remains substantially constant. Thisfacilitates the establishment of steady-state conditions for reductionof the iron contained in the slag to metallic iron. Additionally, rapidcirculation of slag through the oxidising region of the primary smeltingcircuit permits the large amounts of heat which are generated during thehighly exothermic reaction of copper sulphide containing dissolved FeSwith the technically pure oxygen to be transferred to the slag andcarried thereby to the endothermic reduction reactions which are takingplace in the hearths 24 and 28 of the secondary smelting circuit.

In a modification, the apparatus and method of FIGS. 1 and 2 is used tosmelt a mixture of the above-described copper sulphide ore concentrateof high intrinsic energy in admixture with a high grade zinc concentratewhich are mixed together and fed to inlets 32. In this process, zinc isremoved as vapour from R-H vacuum degassing unit 14 and passed to acondenser (not shown) to be recovered as metallic zinc. The zinc isproduced by reaction of copper metal dissolved in the circulating mattewith the zinc sulphide also dissolved in the circulating matte to formcopper sulphide and zinc in an endothermic reaction for which the heatis supplied from the oxidation station in hearth 12 by the circulatingmolten matte. Recovery of iron from the slag is effected in the mannerdescribed above. However, depending upon the energy requirements of thehigh grade zinc ore concentrate and the intrinsic energy of the coppersulphide ore concentrate, recovery of iron from the slag may or may notbe conducted. If it is not conducted, then the secondary smeltingcircuit can be dispensed with.

Typical examples of high intrinsic energy copper are concentrates are:

    __________________________________________________________________________    (1) Inco High Grade Copper Concentrate:-                                             Cu      Ni  Fe      S   SiO.sub.2                                      % wt   30-32   1.1-1.2                                                                           29-30   31-33                                                                             2                                              (2) Kidd Creek Copper Concentrate:-                                                     Cu                                                                              Fe      S Pb    Zn                                                                              SiO.sub.2                                       % wt      26                                                                              28      31                                                                              0.6   4 4                                               (3) Mount Isa Mines Copper Concentrate:-                                      Cu Fe S  Pb  Zn  As  Bi  SiO.sub.2                                                                         Al.sub.2 O.sub.3                                                                  CaO                                                                              MgO                                       24.6                                                                             29.2                                                                             31.9                                                                             0.14                                                                              0.34                                                                              0.18                                                                              0.01                                                                              9.7 1.06                                                                              0.9                                                                              1.13                                      __________________________________________________________________________

The Inco material (1) above has very little gangue oxide associated withthe sulphides, and the slag bleed (line 64) would need to be only about15% of the untreated total slag mass (ie 85% saving in the slag mass tobe disposed of).

Clearly as the gangue oxides in the concentrates fed to the smelterincrease, the relative reduction in mass of slag effected by ironproduction decreases.

Nickel ore concentrates usually have a relatively high gangue oxidecontent compared with copper concentrates, and also the Ni/Fe ratio issubstantially less than the Cu/Fe ratio. Typical examples are:

    ______________________________________                                                        Ni   Cu     Fe     S    SiO.sub.2                             ______________________________________                                        (4) Inco Copper Cliff 11     2    40   30   7                                 (5) WMC Kambalda (Australia)                                                                        13.9   1.1  27   26   17.6                              ______________________________________                                    

Whilst direct production of nickel metal in a single step process ispossible, it is non considered to be economically viable because ofexcessive nickel oxide loss to slag, unless the iron content isextremely low. Thus, with higher iron contents, it is preferred torecover the nickel as refined nickel matte which can then be direct O₂smelted to produce nickel metal.

High grade zinc ore concentrates usually have very low gangue oxidecontents and relatively low iron contents. Typical examples are:

    __________________________________________________________________________    Zn   Pb  Cu Cd As  Fe S  SiO.sub.2                                                                         CaO                                                                              MgO                                                                              Al.sub.2 O.sub.3                           __________________________________________________________________________    (6)                                                                             55.6                                                                             0.03                                                                              0.4                                                                              0.2                                                                              0.2 7.9                                                                              33 0.9 0.25                                                                             0.39                                                                             0.91                                       (7)                                                                             55.0                                                                             1.5 0.3                                                                              0.2                                                                              0.10                                                                              5.0                                                                              32 2.2 1.0                                                                              0.20                                                                             0.48                                       __________________________________________________________________________

Co-smelting of either of (6) or (7) above with a chalcopyriticcopper-concentrate such as (3) above permits 1 tonne Zn and 1 tonne ofcopper to be produced in the melt circulation reactor without externalfuel input if pure technical oxygen is used, provided no iron productionin the secondary circuit is attempted. Assuming that H₂ SO₄ is producedin a single contact acid plant and that the excess O₂ is recycled backto the top blowing region, this leads to virtually zero gas emissionsmelting (some gaseous bleed-off of accumulated nitrogen or air leakageis only necessary).

To produce Zn +Cu+H₂ SO₄ as above requires 16 GJ per 1 tonne zinc plus 1tonne copper compared with 42 GJ per 1 tonne zinc plus 1 tonne of copperfor the best available current technology assuming the overallefficiency of electricity generation is 32%, ie current technologyrequires about 2.6 times more energy/input.

Contact of zinc ore concentrate feed solids with the circulating ironoxide-containing slag should be avoided. Therefore, if metallic ironproduction is envisaged for the zinc/copper co-smelting case, theexternal slag loop material is added to the oxidising hearth 12 (O₂ topblow region) only and then returned to the secondary circuit or externalslag loop without flowing through the other side (ie the first hearth10, see FIG. 1) of the primary smelting circuit.

The R-H unit 14 may be used to effect vigorous contacting between thecirculating matte and the slag recycled from the external slag loop toensure high intensity conditions for heat transference between the slagand matte carrier systems. This can be effected by equipping the R-Hunit 14 with four snorkels as opposed to the two (as illustrated in FIG.1). Two of these would be for circulating matte, whilst two would be forslag. They could be arranged so that one or both phases are dispersedthrough the other to promote very high intensity heat transfer.

Instead of using the R-H unit 14 in this way, the top blow region itselfcan be arranged to promote efficient heat transfer between the carrierslag phase and the underlying matte phase. Attention must be paid to theslag clearing propensity of the top blow lances 40 so that a free mattesurface is available under the impact. zones of the gas jets foreffective reaction between O₂ and matte. Care will need to be exercisedin this area if Zn or Pb smelting is occurring along with copper. Themass transfer of zinc sulphide dissolved in matte from the bulk to thetop surface of the matte must be restricted and so it is consideredadvisable to effect relatively soft top blowing in the non-splashregion.

It will usually be necessary in all the above-described schemes toeffect in situ slag cleaning as the slag floats away from the top blowregion by adding pyrite for example to the slag layer and then allowingit to settle out. However, the external slag loop provides additionalsafety. An intermediate iron carbon melt containing dissolved Cu and Pbin particular can be bled off and returned (see line 54--FIG. 1) to theprimary matte circuit where the reaction Fe+Cu₂ S→FeS+2 Cu takes placegenerating copper whilst any copper originally in the slag and now inthe iron bleed stream is recycled back to the main copper smeltingfurnace,

I claim:
 1. A method of oxygen smelting of a sulfide ore concentrate ofhigh intrinsic energy value selected from the group consisting of acopper sulfide ore concentrate of high intrinsic energy value, a nickelsulfide ore concentrate of high intrinsic energy value and a copper andnickel sulfide ore concentrate of high intrinsic energy value,comprising the steps of:(a) forcibly circulating a molten sulfidecarrier composition through a closed loop extraction circuit from whichat least one product selected from copper, nickel and sulfides thereofcan be continuously extracted at an elevated temperature; (b)introducing said sulfide ore concentrate of high intrinsic energy valueinto said molten carrier composition at an ore receiving station so thatsaid sulfide ore concentrate of high intrinsic energy value is dissolvedin or melted by said molten carrier composition; (c) contacting saidmolten carrier composition containing said sulfide ore concentrate ofhigh intrinsic energy value with an oxidizing gas containing at least 30volume % oxygen at an oxidation station so as to oxidize at least partof said molten carrier composition containing said sulfide oreconcentrate of high intrinsic energy value; and (d) utilizing heatgenerated during the oxidation step as a result of oxidation of saidsulfide ore concentrate of high intrinsic energy value by smelting saidsulfide ore concentrate of high intrinsic energy value with anothermineral sulfide concentrate of low or negative intrinsic energy valueselected from the group consisting of high-grade zinc sulfideconcentrate, high-grade lead sulfide concentrate and a bulk flotationsulfide concentrate containing both lead and zinc.
 2. The methodaccording to claim 1, wherein said heat utilizing step (d) alsocomprises reducing iron oxide in slag produced in the method to a liquidiron product; and wherein said liquid iron product is recovered.
 3. Amethod of oxygen smelting of a sulfide ore concentrate of high intrinsicenergy value selected from the group consisting of a copper sulfide oreconcentrate of high intrinsic energy value, a nickel sulfide oreconcentrate of high intrinsic energy value and a copper and nickelsulfide ore concentrate of high intrinsic energy value, comprising thesteps of:(a) forcibly circulating a molten sulfide carrier compositionthrough a closed loop extraction circuit from which at least one productselected from copper, nickel and sulfides thereof can be continuouslyextracted at an elevated temperature; (b) introducing said sulfide oreconcentrate of high intrinsic energy value into said molten carriercomposition at an ore receiving station so that said sulfide oreconcentrate of high intrinsic energy value is dissolved in or melted bysaid molten carrier composition; (c) contacting said molten carriercomposition containing said sulfide ore concentrate of high intrinsicenergy value with an oxidizing gas containing at least 30 volume %oxygen at an oxidation station so as to oxidize at least part of saidmolten carrier composition containing said sulfide ore concentrate ofhigh intrinsic energy value; (d) utilizing heat generated during theoxidation step as a result of oxidation of said sulfide ore concentrateof high intrinsic energy value by reducing iron oxide in slag producedin the method to a liquid iron product; and (e) recovering said liquidiron product.
 4. The method according to claim 1, wherein said oxidizinggas is technically pure oxygen.
 5. The method according to claim 3,wherein said oxidizing gas is technically pure oxygen.
 6. The methodaccording to claim 1, wherein copper is extracted as copper metal andnickel is extracted as nickel sulfide.
 7. The method according to claim3, wherein copper is extracted as copper metal and nickel is extractedas nickel sulfide.
 8. The method according to claim 1, wherein saidsulfide ore concentrate is exposed to radiant heat before it isintroduced into said molten carrier composition so as to vaporize labilesulphur and other volatile sulfides, and wherein said vapors are removedand condensed.
 9. The method according to claim 3, wherein said sulfideore concentrate is exposed to radiant heat before it is introduced intosaid molten carrier composition so as to vaporize labile sulphur andother volatile sulfides, and wherein such vapors are removed andcondensed.
 10. The method according to claim 8, wherein said sulfide oreconcentrate is exposed to radiant heat by causing it to pass downinclined walls at said ore receiving station before introduction intosaid molten carrier composition.
 11. The method according to claim 9,wherein said sulfide ore concentrate is exposed to radiant heat bycausing it to pass down inclined walls at said ore receiving stationbefore introduction into said molten carrier composition.
 12. The methodaccording to claim 3, wherein the heat utilizing step (d) and the liquidiron product recovering step (e) are conducted in a secondary circuitutilizing a molten slag phase removed from said closed loop extractioncircuit.
 13. The method according to claim 12, wherein (i) reduction ofsaid iron oxide is effected by addition of reductant to said secondarycircuit, (ii) the amount of reductant added to an upstream part of saidsecondary circuit is restricted so that hot metal initially formedcollects residual copper and other reducible element impurities thatleave said closed loop circuit dissolved in the slag, and (iii) theinitially formed hot metal is collected and returned to said closed loopcircuit.